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By Frederick Bird

This paper is based upon a review of studies of mostly Canadian mining companies in Chile, northern Canada, Tanzania, Guatemala, Ghana, Sierra Leone, and the Democratic Republic of Congo. In spite of often well-meaning efforts most mining firms produce wealth in developing areas that largely benefit those immediately involved, sometimes neighboring communities, and often those in the governing strata. Typically mining takes place in enclaves and fosters enclave development, rather than the kind of inclusive wealth development favored by the recently UN published Inclusive Wealth Report. One way by which mining operations can foster more inclusive wealth development is to adopt much broader approaches towards the construction of the various physical and social infrastructures ? roads, water, electricity, health services, schools, security ? required for their own operations by collaborating with diverse regional, provincial, and/or national agencies and businesses to develop these infrastructures for larger populations and areas beyond the immediate sites where they are operating. Established mining firms can also foster more inclusive wealth development by establishing more formal, collaborative relationships with artisanal miners.

Aug 1, 2013

By S. Tuttle, R. Sisson

"Suncor Energy Inc. has proposed a 2.2 billion dollar expansion to the current oil sands operation, which will more than double the productive capability of the facility. As part of the application for this expansion, a comprehensive closure plan had to be developed and filed.Closure planning includes a systematic evaluation of the area to be developed, the development activities planned, the goals and objectives of the corporation in re-establishing landforms and ecosystems at the same time as running an effective operation. As early as is practical during the mine development, surface contouring is completed, soil is reconstructed, and vegetation, surface drainage and wetlands are re-established. Once active reclamation is complete and vegetation has been re-established, Suncor monitors maturation of landscapes and ecosystems to provide the submission basis for reclamation certification"

Jan 1, 1999

By I. P. Barbosa, P. S. M. Soares, I. O. C. Masson

"Permanent magnets have been applied extensively to several electronic products such as printers, TV sets, microphones, microcomputers and headphones, presently manufactured in large scale by industrial economies. Sm-Co alloys represent about 20% of the world market for permanent magnets and their industrial processors generate 20-30% of magnetic scraps which can be recycled. This paper discusses a process route to recycle SmCo5 using acid treatment, solvent extraction of Sm and electrowinning of Co. The results confirm the feasibility of the alternative proposed. Basic operational variables are suggested as guidelines to design a plant for recycling magnetic scraps.The use of hydrometallurgical processes to recycle scraps from the industry of permanent magnets has been suggested in some recent papers related to recycling of metallic alloys [ 1 ]. These research efforts might be understood as a worldwide tendency towards the development of self sustainable metallurgical processes which mean the reduction of environmental impacts.CETEM -Center for Mineral Technology leads in Brazil several research programs concerning recycling and treatment of industrial residues. The work on samarium purification and concentration by solvent extraction, producing also a solution of cobalt sulfate as partially reported here is one of these programs. Previous results related to the sulfuric acid digestion of SmCo5 are presented somewhere else [2]"

Jan 1, 1997

By Steve Cole

"The economic and strategic importance of PGE bearing deposits is fuelling renewed exploration and development efforts in the precious metals mining sector.This paper presents a review of the various techniques used in the processing of PGE bearing ores, from primary benificiation to processing routes for the different flowsheets that have been developed amongst current producers. Recent innovations such as preconcentration of ore using gravity separation, recent trends in flotation circuit design and the development of hydrometallurgical routes for the direct treatment of flotation concentrates without the need for smelting will be discussed. Data from recent projects will be presented to illustrate common characteristics of PGE ores and concentrates.INTRODUCTIONThe six platinum-group elements (PGE): ruthenium, rhodium, palladium, osmium, iridium and platinum, together with gold and silver, are considered to be ""precious"" metals. The recent spike in platinum and palladium prices coupled with the projected increase in demand for the metals has fuelled a spate of exploration and new projects in the PGE sector. Historically primary producers of PGE metals have controlled the flow of information relating to the mining and processing of these ores into the public domain. Very little published information relating to individual operations and processes is available in comparison to the abundant literature relating to other metals.The purpose of this paper is to provide a general overview of the issues relating to the processing of PGE ores. Each topic discussed could easily be the topic of a paper or book and so in many cases only the surface of the topic will be scratched. An attempt has therefore been made to provide as many references as possible.This paper discusses the classification and location of PGE ores, the mineralogical factors that affect recovery of PGE's, methods currently used for recovering PGE's and process trends in PGE metallurgy."

Jan 1, 2002

By Andrew G. Patton

The Alimak mining method was introduced at the David Bell mine in 1998 to extract an otherwise uneconomic and inaccessible reserve. The success associated with the extraction of these reserves led to major expansion of the method. To date, results have been extremely positive in terms of grade control, reduced dilution, stope recoveries, productivities and mining costs. It is hoped that through further development of the method at David Bell mine we will continue to see positive results in these areas and be able to increase mineable reserves to include proven resources that do not currently meet the criteria of our current stoping plan.

Oct 1, 2001

By David Seaman, J-P Franzidis, Emmy Manlapig

"A new device has been developed to directly measure the bubble loading of particle-bubble aggregates in industrial flotation machines, both mechanical flotation cells as well as flotation column cells. The bubble loading of aggregates allows for in-depth analysis of the operating performance of a flotation machine in terms of both pulp/collection zone and froth zone performance. This paper presents the methodology along with an example showing the excellent reproducibility of the device and an analysis of different operating conditions of the device itself.The device has been used to investigate the froth selectivity of attached particles. Simultaneous measurements were made using a different froth recovery measurement technique in the same flotation cell. A comparison of these techniques shows that in this case, particles are selectively detached from bubbles at or near the pulp-froth interface.INTRODUCTIONA flotation cell can be considered to operate in two distinct phases or zones; namely the pulp or collection phase and the froth phase. The pulp zone is responsible for the formation of hydrophobic particle-bubble aggregates, whilst the froth phase is responsible for the separation of these aggregates from surrounding suspended hydrophobic and hydrophilic particles. Thus, the froth phase is the grade determining phase in flotation. A measurement of the bubble load (the mass of particles attached to air bubbles in units of mass/volume) will enable the analysis of the performance of both the collection zone at creating bubble-particle aggregates and also the froth zone at separating these aggregates from surrounding pulp.A measure of froth zone performance that has been used extensively in flotation research is the Froth Recovery, Rf, of a particular system. Froth recovery is defined as the mass of particles collected in the concentrate launder by true flotation as a fraction of the mass of particles entering the froth phase. The measurement of bubble load allows for this term to be calculated easily, rather than having to go through the laborious task of, for example, changing froth depths to determine this performance criterion (Contini et al., 1988; Vera et al., 1999b; Feteris et al., 1987)."

Jan 1, 2004

By L. Lui, A. Cauchon, A. R. Laplante

"The gravity circuit at Les Mines Camchib Inc. consists of two pinched sluices to bleed a feed to one of two 76 cm (30"") Knelson concentrator. The Knelson concentrate is upgraded to smeltable grade with a smaller Knelson (18/30 cm) and a riffleless table. Their tails are column cyanided and washed prior to recycle to the grinding circuit. This paper details size-by-size circuit performance. The Knelsons recover around 90% of free gold (as determined by Mozley Laboratory Separator), even below 37 µm. The sluices are ineffective, with upgrading ratios around 2, but approaching unity in the fine sizes. Much of the effectiveness of the circuit lies in the high gold circulating loads in the grinding circuit, which partly atone for the poor sluice performance, and the ability of the Knelsons to recover free gold. Secondary upgrading recovers slightly above 99% of the primary concentrate gold, 90% of it in a 65% Au concentrate and the balance in a heavily loaded (10 to 20 oz/st) pregnant solution. INTRODUCTIONBackgroundThe operating performance of the mill of Les Mines Camchib Inc. has been described earlier (Laplante et al., 1989). The presence of significant chalcopyrite (1 % and up) makes flotation necessary prior to cyanidation, to limit cyanide consumption and minimize environmental impact. In 1982, the presence of very high gold circulating loads in the grinding circuit lead to the commissioning of a gravity circuit, schematically shown in Figure 1. Gravity recovery of gold has been found beneficial because it increases metallurgical recovery by 1-3% (Laplante et al, 1987), as only part of the flotation tail is cyanided, and economic recovery, because of the lower net smelter return of gold in the flotation concentrate (as compared with the gravity concentrate). The latter increase is estimated, with a net smelter return of 95%, at 1.5% when gold recovery by gravity is equal to 30%. In early 1988, a sampling program was initiated to characterize the behaviour of gold in the grinding/gravity circuit. This paper reports on the results of this study."

Jan 1, 1990

By Nick Poushinsky

OVERVIEW OF PRESENTATION ?Stantec?s Northern Presence ?Challenges for Northern Development ?Environmental Approval Challenges ?History/Expectations ?Baseline Data ?Consultation ?Regulatory Processes ?Extra ?Regulatory ?Keys to Success

May 1, 2009

Optimizing production scheduling in open pit mines is important for managing cash flows of mining operations. This paper describes a new, probabilistic, mixed integer programming (MIP) formulation which has been developed to minimize the risk of not achieving planned production targets, in terms of metal production and ore tonnage feed for the mill during production, while considering maximization of total discounted economic value. The new MIP model also recognizes that to minimize movement of excavators over long distances and accommodate equipment access, the blocks scheduled for mining within a period cannot be spread too widely. The model is applied to a laterite nickel-cobalt deposit, as an example of multi-element deposits that have complex ore quality constraints. The results show that the proposed model generates feasible schedules. When compared with traditional optimization techniques, the new model is superior in terms of minimizing the risk of not meeting production targets, especially at the early stages of production, and in terms of generating practical scheduling patterns.

Jan 1, 2006

"This paper reviews the chemistry involved in the recovery of gold from cyanide solution with anion-exchange resins, and also discusses the most critical aspects of the engineering of an effective resin-in-pulp (RIP) process. As far as the chemistry is concerned, the conclusions reached by a great many researchers, as well as the performance of the only operating RIP plant in the Western World, at the Golden Jubilee Mine in South Africa, support the view that the process is technically viable and, in many cases, could prove to be more attractive economically than alternative processes such as zinc cementation and carbon-in-pulp (CIP). On the engineering side, there is still a need for further development before larger scale ( > 50,000 tpm) RIP plants can be designed, built and operated with confidence. However, the experiences gained during nearly 3 years of operation of the Golden Jublee RIP plant indicate that the process can be engineered with confidence for smaller operations ( < 50,000 tpm), and this will probably be the scope of application of the first-generation RIP plants. INTRODUCTIONThe use of anion-exchange resins in gold processing has, for a number of years, been proposed as an alternative to more conventional processes such as zinc cementation and carbon adsorption, and detailed research programs that have been carried out at a number of laboratories have demonstrated the technical viability of the process. (Burstall et al, 1953, Davison et al., 1960, Tataru, 1973, Laskorin et al. 1978, Hosking, 1984, Riveros and Cooper, 1987 and Palmer et al., 1988)The main incentive for this work was the fact that resins have several characteristics that are quite different from carbon, which make them potentially more attractive substrates for the recovery of gold. The following are examples: -"

Jan 1, 1991

By J. E. Sepulveda, R. Morrow

"In full scale mineral grinding operations, the steel balls normally utilized as auxiliary grinding media get to be consumed at different specific rates, as determined by 3 basic mechanisms: abrasion/corrosion (wear), impact ball breakage and impact surface spalling, the latter relating to the removal of rather small pieces of steel coming off the exposed surface of the balls due to repetitive impact with other balls in the charge or the surrounding mill liners. In conventional ball milling applications, the clearly predominant mechanism is gradual wear by abrasion/corrosion. In the case of SAG mills, where the make-up balls may be as large as 6”f diameter, breakage and spalling mechanisms are significant contributors to total ball consumption, although not to the extent of overriding wear mechanisms.Ball wear has been well characterized by the Linear Wear Theory, as presented at the last CMP 2014. Impact breakage has been customarily monitored, at the ball manufacturing site, via Drop Ball Tests (DBT) consisting of repeatedly dropping balls from a standard height and accounting for the number of broken balls and the weight loss (spalling) of the remaining unbroken balls. When properly conducted, DBT’s should provide meaningful indications of the breakage and spalling resistance of large diameter balls. The current publication is focused on the interpretation and projection of DBT results to full industrial scale, taking into account all 3 mechanisms of grinding media consumption in SAG mills.INTRODUCTIONA thorough, quantitative understanding of the independent contribution of all active grinding media consumption mechanisms is of the highest practical relevance, particularly in semiautogenous (SAG) milling operations where ball breakage and spalling can not be ignored or neglected, as it has been normally acceptable for conventional ball milling processes. The high severity impact conditions currently imposed in typical SAG mill environments create serious challenges to the manufacturers and suppliers of large diameter (5”–6”) SAG balls, whose products must be capable of surviving such harsh operational conditions with minimum breakage or spalling."

Jan 1, 2015

By Quin Dennis

Process Optimization ?What is it really? It is NOT ?a larger fan to increase draft through a process ?a larger pump on a wet circuit ?a wider conveyor belt for solid materials ?a larger feed elevator into or out of a process These relieve individual constraints ??bottlenecks??and allow operation to the next constraint in the process ?It is also NOT about automating the operator ?which captures consistency but is not continuous optimization

Nov 1, 2011

By Ronald L. Bruening

IBC has developed a cost effective MRT process for the selective extraction and recovery of gold, copper, and a mixture of other base metals from mine leach cyanide solutions. This is a two-stage system in which the gold is first recovered selectively as a high purity salt. The copper is then recovered separately as a sulfate salt. The other base metals are then collected as a cyano anion concentrate. Three different SuperLig® materials are used. The first stage can be used alone on gold leach solutions where copper and other base metals are not a problem. The two or three stage system is used when the base metals build up in the leach solution, resulting in increased costs. The cyanide or cyanide/cyano anion concentrate can be recycled back to the leach solution until the build-up of other salts, etc. becomes too high, or this too can be controlled with a percentage bleed stream.

Oct 1, 2003

By P. Dyas, H. Zouitt, D. M. Smith

"Falconbridge Copper's Lac Shortt gold mine is located in northern Quebec approximately 300 km northeast of Val D'Or and 100 km west of Chibougamau (Pig. 1). Engineering for the mill commenced in early 1983 with a feasibility study. The plant was commissioned in the fall of 1984 at a capital cost of approximately $12 million. Design capacity was 750 tpd which was increased in the fall of 1985 to the current level of 1200 tpd.The ore at Lac Shortt is free milling, grading an average of 5.5 g/t gold which is finely disseminated throughout the host rock. Pyrite at 3 to 5% by volume and traces of copper mineralization are also present.Some of the latest gold processing techniques known at the time of design were incorporated into the mill. These include carbon-in-pulp, pressure stripping, electrowinning of pregnant solution and induction furnace refining. The plant also features several innovations such as a unique interstage carbon retention screen in the CIP circuit and the outdoor location of the cyanidation feed thickener, mill solution tank and reagent storage silos despite -40° Celsius winter temperatures.This paper describes the current mill process and discusses some of the problems encountered during the first two years of the plant's operation. CRUSHINGThe Lac Shortt surface crushing facilities are located in a separate building, isolated from the wet processing circuits for dust and noise control reasons. The crushing plant equipment was originally utilized at the Manibridge operation in northern Manitoba where it had been designed to operate at 135 tph. However, at Lac Shortt a rate of only 80 tph has been achievable to date, mainly due to the hardness and frequently slabby breakage characteristics of the ore. Therefore, in order to meet fine ore requirements at the current mill throughput of 1200 tpd, the crushing circuit is run seven days a week. 24 hours a day. Availability has been averaging 65%."

Jan 1, 1987

By Brad Thrall

"Viceroy Resource Corporation of Vancouver, B.C. has recently developed the Brewery Creek Mine, an open pit heap leach gold mine 55km East of Dawson City in Yukon, Canada. Mineable reserves are 17.1 million tonnes of ore grading 1.45 grams/tonne, containing 24.8 tonnes (797000oz) of gold. Ore processing employs the heap leach technology using run of mine ore. 1996 highlights included construction of 158,000 m2 of leach pad, placing 1.9 million tonnes of ore at a grade of 1.46 gpt on to the leach pad, construction and successful commissioning of the ADR plant, and pouring of the first gold bar on November 15th.Permitting this mine has drawn significant attention because the heap leach technology, while proven in other jurisdictions, has not been used in Yukon where winter site temperatures routinely reach -45 C for extended periods.IntroductionFollowing a comprehensive governmental and public review of the proposed project, a Class A Yukon Water license was issued in August, 1995. Construction started in 1995 and leaching commenced in October 1996. The Brewery Creek Mine is in Yukon, the home of the Klondike Gold Rush nearly 100 years ago. This is the only part of Canada not glaciated in the Wisconsin advance so there is oxide ore resulting from about 80,000 years of weathering -unique in Canada. Discontinuous permafrost is present on site and has required special consideration in the design. Winter temperatures dip below minus 50 C, demanding an approach to mining and leaching that is different than in more temperate climates.Project DescriptionThe mine comprises 8 shallow, largely oxidized gold deposits along a 7km corridor. An 8 year mining program will extract 17.1 million tonnes of ore grading 1.45 grams/tonne, containing 24.8 tonnes (797,000 oz) of gold. The strip ratio will be 1.35: 1. Overall gold recovery is projected at 78%. There will be 2.3 million tonnes of ore placed on the pad annually, with an average of 2.57 tonnes (83, 000 oz) of gold recovered per year. Because of the cold winters, the mine will operate approximately 210 days per year, while the leaching process will operate on a year round basis. Run of mine ore is placed on the heap 150 days per year. The recovery process utilizes conventional sodium cyanide leaching , carbon adsorption, caustic desorption, and on-site smelting. Peak employment during mining is 145 and winter staff is reduced to 45."

Jan 1, 1997

"Gravity concentration remains the main separation method for fine iron ore and is used extensively for treating various minerals. Gravity concentration methods separate minerals of different specific gravity by their relative movement in response to gravity and one or more other forces, the latter often being the resistance to motion offered by a viscous fluid, such as water or air (Wills, 2006).Spirals concentrators are one of the gravity concentration equipment used for beneficiation of minerals. Spirals are typically used for processing of fine minerals (-1mm) size fraction. The main operating variables for a spiral concentrator are the feed rate, the solids concentration, the wash water addition and the position of the cutters or splitters used to separate the tailings, middling and concentrate streams (Sadeghi, Bazin & Renaud, 2014).The aim of this study is to investigate the effect of feed rate, solids concentration and wash water on grade and recovery on iron ore material. The Multotec SC20HC/7 WW spiral was used for the experimental test work. The spiral was tested at varying solids concentration and feed rate. The wash water rate was also varied to establish the best washing rate. The splitter positions were kept constant. Optimization of these variables is expected to provide optimum conditions for a good compromise between grade and recovery.INTRODUCTIONSpiral concentrators are simple low energy consuming devices that separate minerals mainly on the basis of density. Spirals are widely used in mineral processing as a method for preconcentration and have proven to be metallurgically efficient and cost effective.Spirals have a wide application, they are used in iron ore processing, coal, gold, chromites, mineral sands, glass sands, and in soil cleaning.Separation on a spiral is achieved through a combination of forces that act on particles as they move down the trough of the spiral. The main forces known to act on the particle on a spiral are the gravitational forces, centrifugal force, hydrodynamic drag, and lift and friction forces (Kapur & Meloy, 1998). Apart from the forces acting on a spiral, the properties of the slurry flowing on a spiral including, solids concentration, feed rate and wash water also plays an important role in the separation on the spiral.The relationship between separation efficiency and feed rate was shown to be linear, with a decrease in efficiency at higher flow rates. Previous work on flow profile by Holland - Batt (1990) indicated that any increase in the feed rate to a spiral results in a greater part of the additional flow volume reporting to the outside of the trough. The flow on the outside of the trough becomes energized and consequence to that the heavy particles find it hard to settle and escape this region to the inner region. This effect has a high impact on recovery of heavy minerals, since any increase in volumetric flow on the spiral is not made available for recovery in the inner region of the spiral."

Jan 1, 2015

By Ralph Harris, Qi Xia

"Gold cyanidation rates are known to be limited by the rate of oxygen dissolution in the cyanidation tanks. The effect of solids loading on the oxygen dissolution rates is not obvious and the results of previous studies are ambiguous. The present paper presents an analysis of oxygen dissolution rates using a ""spent"" commercial gold slurry diluted to various levels of solids loading. It was found, in an 170 litre stirred tank reactor using an inverted cone sparger, that under certain high loading conditions, there was a decrease of as much as 50 %, i.e., there was a halving of the rate of oxygen dissolution as compared to that in clean water. In particular, it was observed that, due to gas flooding of the impeller, there was a drop in kLa even with an increase of gas velocity. Gas holdup reached a maximum value of only 1.7 %. The total power consumption was between 0.06- 0 .10 kw/m3 within the range of gas velocity before flooding occurred. The decrease in oxygen dissolution rate with increased solids loadings is believed to be mainly due to the increase in the apparent viscosity which reduced the turbulence and shearing effect and resulted in larger bubbles and hence reduced the reaction area a. The interrelations between solids loading, gas holdup, flow pattern, power consumption, and liquid mass transfer coefficient are discussed.IntroductionThe reactors generally used in gold cyanidation are Stirred Tank Reactors (STR). Compared with other types of reactors like fluidized bed reactor and column reactor, STR is of particular importance due to its wide application in industry. It is reported that the stirred tank reactor is used for more than 50 % of the world's chemical production (Xu and McGrath, 1996)."

Jan 1, 1997

By W. H. MacLean, S. Liaghat

Abstract- The Key Tuffite is a thick (0.3 m to 6 m) and laterally extensive (-10 km) cherty, sulfide-bearing tuff unit in Archean greenstones in the Matagami mining district in northwestern Quebec. It lies at the contact of the Watson Lake Group and overlying Wabassee Group volcanic rocks, and is host to major Zn-Cu volcanogenic massive sulfide deposits. The tuff component is highly altered and its volcanic precursors are not identifiable by petrographie means. The tuff is mixed with a volcanic exhalative component comprising chert, carbonate and sulfides (pyrite, pyrrhotite, and minor sphalerite and chalcopyrite). The normative mineral composition of the Key Tuffite is 33% quartz, 28% sulfides, 16% sericite, 8% carbonate, 60Jo chlorite and 9% other minerals. In the vicinity of the Bell Allard and Orchan mines, the precursor tuff components are identified using immobile major and trace elements, as Watson Lake tholeiitic rhyolite (Zr/Y = 1.5 to 3.5, LaN/Ybn. = 1.7 to 2.8, Si02 = 72.5 ± 1.2 wt %, and Wabassee calc-alkaline andesite (Zr/Y =9, LaN/Ybn. = 4 to 8, Si02 = 56% to 65%). Immobile element computation techniques were used to calculate mass additions and losses in the conversion of the tuff to the Key Tuffite. Average net amounts of Fep3 (27% total iron), Si02 (14%), CaO (8%), S and K20 ( + 2%), were added, Na2O was removed (-2.8%), and MgO was over-all unchanged. Total net mass change was +63%, which comprises mainly exhalative additions, but also includes changes owing to hydrothermal alteration of the tuff component.

Jan 1, 1992

By Paul Harvey

Following initial discovery in 1991, BHP Billiton Diamonds Inc opened Canada?s first diamond mine, the Ekati Diamond Mine TM (Ekati) in the Northwest Territories on October 14th, 1998. Four short years later on November 29th, 2002, saw the opening of North America?s first underground diamond mine - the Koala North mine. Nestled between the original Panda open pit and the new Koala pit which is in the early stages of production, Koala North represents the only underground kimberlite mining venture in arctic conditions outside the new underground mines being developed by ALROSA in the Yakutsk region of eastern Russia. Extensive underground kimberlite mining has been practiced in the Republic of South Africa for much of the 20th century and continues today. Underground mining represents a key component of the long term business plan for the Ekati. As various open pits reach their economic limit, continued recovery of high value ore at depth is dependent upon the ability to productively mine the kimberlite pipes by underground methods. As a function of its location, size, and value, the Koala North pipe is ideally positioned to host the first underground mine at Ekati . In addition to being a production source for the operation, Koala North is designed as a learning mine, from which to gain valuable experience in preparation for the potential mining of other larger, higher value pipes by underground methods in the future. Fortuitously Koala North was available for development so the learnings could be integrated into the Panda mine design prior to completion of the Panda pit. Establishing our first underground diamond mine, and applying the open benching mining method in arctic conditions brings a unique combination of mining challenges together in a harsh climate and remote environment. This paper examines the design and establishment of the Koala North mine. Discussion includes experiences gained from the first year of development, production start-up, and provides a brief overview of our potential underground future at Ekati .

May 1, 2003

By P. Wells

"In the past decade, there have been many developments in processing of nickel/copper sulfide ores. Improved pentlandite/pyrrhotite flotation selectivity with the use of new reagents, fine regrinding and Redox measurement (control) has led to increased pyrrhotite rejection and thus higher concentrate grades. Flash flotation has increased overall nickel and associated precious group metal recoveries. Recovery of both fine and coarse particles has been improved through an understanding of flotation cell hydrodynamics. The separation of copper from nickel in such ores has seen a similar level of development. From the late 80s to very recently, the use of flotation columns, the use of pyrrhotite depressing reagents and the improved understanding of the complex mineralogy of chalcopyrite, pentlandite and pyrrhotite have considerably increased copper / nickel separation efficiency. Copper concentrate with more than 30% copper and less than 0.5% nickel grade is now obtainable at 80-85% copper recovery from mill feed compared to 70-75% copper recovery with 30% copper and 1% nickel grade only a few years ago. Improved copper / nickel separation in milling operations has the potential to substantially improve the processing of the resulting concentrates and thus the overall return from the extraction of nickel and copper from sulfide ores.INTRODUCTIONNickel metal by itself or in alloys with other metals and materials makes a significant contribution to modern society. The relatively high price that nickel commands compared to other base metals is a reflection of the fact that, although abundant in the universe, nickel is sparingly distributed in the earth’s crust. There are two types of nickel deposits that are currently being exploited: complex sulfide ores and lateritic oxide ores. Lateritic contain no copper but in sulfide deposits, nickel– bearing minerals, principally pentlandite (Pn, [Ni,Fe]8S9), some millerite (NiS) and violarite ([Ni,Fe]S) are often found in association with copper-bearing minerals, mainly chalcopyrite (Cp, CuFeS2) and cubanite (CuFe2S3). The copper content of such ores can vary considerably from as high as 5-8% (in some Sudbury basin ores) to as low as 0.1% (in ores mined by Western Mining at Mount Keith, Western Australia and by Inco in Thompson, Manitoba). The valuable minerals Pn and Cp are almost invariably found associated with pyrrhotite (Po, Fe7S8) which along with nonsulphide rock has to be rejected to tails in the primary milling stage. The bulk concentrate from the primary stage is then treated to depress Pn and float Cp. A complete Cu/Ni separation at the mill would be highly desirable but this is not possible because of locking between a small fraction of the valuable minerals. As a result, Cu/Ni separation has to be carried out both during milling and during refining of the matte following smelting. At Inco, Jinchuan and in Norilsk, a chalcocite / heazlewoodite separation is made by flotation following slow cooling of the matte. In the present Sudbury Inco operations, Cu/Ni separation has not been conducted at the mill since 1991 and all Cu/Ni separation takes place at the matte processing plant. Because of the rising ratio of Cu to Ni in the ore, this strategy is under review and this paper describes some recent studies of circuits in preparation for possible installation of a copper / nickel separation circuit at Inco’s Clarabelle Mill in the Sudbury area."

Jan 1, 2004